Recovery of metal values from complex ores



o. F. MAr'aI/IN l 3,230,071 RECOVERY OF- METAL VALUES `FOM COMPLEX DRES Fneduay 25. 1952 Jan. 18, 1966 5 Sheets-Sheet 1 oke/N E MMV/A1Y AIToRNEY Jan. 18, 1966 F. MARVIN RECOVERY OF METAL VALUES FROM COMLEX CRES Filed May '25. 1962 5 Sheets`-Sheet 2 ATTURNEY' Jan. i8, 1966 Filed May 25, 1962 o. F. MAR-VIN RECOVERY oF METAL VALUES FRQM COMPLEX oREs v ATTURNEY Jan. 18, 1966 REcovEm oF 'METAL VALUES non COMPLEX oREs Filed may 25, 1962 5 Sheets-Sheet 4.

I oem/v A MeV/N Arroz/wav I NVENTOR.

Jan. 1s, 1966 o. F. MARVIN aEcovrsRY oF METAL vALuEs FROM COMPLEX oRE's Filed may 25, 1962 5 Sheets-Sheet 5 172? f7; Low TEMP 4BULK (6mm 4001?) courir-ruff mvneouf comu Ex l( 6 Amos/HERE oef w/rl/ ff, fj

20A sr Pa, cu ,f zu A5 1N F1a l swr/056 173/ /71/ m5? lef? 1842 ,4c/D misma/E RES/nue Res/DUE RES/Duc LEA cH RoAsrED my TED EAC/QED Afrfk Aff nc To (clem 55a?) mm conc l wir/1' Ac/o 54m efMm/E murmura M2 C05l ACE rfc RoAsrEo H20/V wmp/i521 sou/170A' Acro {tv/RCA coa 51N F1a l HUE/25o oLur/ou cour/ouin AmvsPHEpe Y A5 IN FIG l FEB/ous schuf/0N `ovLunroN v SULFATE )#74 (79 L 50B/UH LEADv '7u/8? 50u/HON .SULPHA rE AmTE I9/ 2 v cu l 1 mec/Pl mrt WITH 'Zu RON Eno y l PEL-DVERY 7*/ 7C `RECOVERY .1 /83 89? CU ELECTRO nsposlrfn 1882 19a sci/T -f lzgmv /9 7" COPPER soLunoN f .efcovfkv v 19s? 19a 1932 192? ma] as??` nfs/DUE Ac/o zealous 25,005 2:3205 COFFEE 0MM/N5 EACH ro eoAsreD O/.o cL RECOVER (cnamA cso'o calm/MN@ LEACHED suLFArE son: z/Nc As 1/mamma zmc @rio saLurloN SILVER THE Arf/esmas@ .Renova suLFA TE A: m F1a; vcoPlsR 199) zal 2 202) es ',REEUDE @om 5f# @om p MTH VER CYAN. su VER awr/o lfzfscove'm' NAN/oe 5 N .ioL ur/oN .VNTOR 1N l Y ofzR/N F. MAW/N B nfs/DUE l 203 ATTORNEY' Unite Stte 3,230,071 RECOVERY F METAL VALUES FROM COMPLEX CRES Orrin F. Marvin, P.0. Box 1035, Cottonwood, Ariz; Filed May 25, 1962, Ser. No. 197,703

8 Claims. (Cl. 'I5-2) My invention relates in general to the recovery of metal values from complex ores. It relates more in particular to improvements in recovery methods as heretofore known Y by means of which the complex minerals of the ore may be broken and the ore becomes available for treatment by substantially any known chemical and/ or metallurgical method to obtain much better recoveries of the totall *will have a deleterious effect on subsequent recoveries,

leaching soluble metal val-ues, thenvheating at a high ,temperature to vsolubilize the 'remaining metal values,

leaching the same, and finally treating the solutions by hydrometallurgical methods to recover substantially all of the metal values.

In my co-pending application S.N. 10,210, iiled February 23, 1960, I disclosed a new process wherein the lirst low temperature roast was then subiected to a reducing roast after which substantially all of the iron but none of the copper was solubilized, leaching the roasted concentrate to remove iron, roasting the residue at a higher temperature, and then leaching out all of the remaining.

metal values and then recovering the same by chemical and /or hydrometallurgical means.

In accordance with each of my prior inventions identified hereinabove, there is a partial breakdown of the complex such that metal recovery is markedly improved, as contrasted with other available methods. One of the limitations of such methods, however, 'is the fact that the recoveries associated with the roasting step or steps are limited in that, at .least to some extent, and particularly with respect to the removal of iron, hydrome'tallurgical methods are employed; and while there is no objection to removal of iron by leaching, and in fact certain definite advantages are obtainable thereby, ank over-all process may indicate a preferable step at this point and the process as a whole shows definitely that there are some limitations inthe re-arrangement of the chemical and/or crystalline bonds characteristicof complex ores..

There is still another aspect of my prior inventions, and that is that they are primarily directed to and have their greatest utility under circumstances when several metals are present in fair concentrations. Most characteristic of this kind of condition is the type of ore in which there is a substantial amount of iron, copper, zinc and lead sulfide with smaller proportions and traces of possibly several other metals including rare metals, silver and gold. But complex ores are by no means limited to the well-known iron-copper-zinc-lead sulfide combinations characteristic of many Western State deposits. Even the relatively simple iron-copper sulfides, of which chalcopyrite is illustrative, and which are mined extensively for their copper content, are actually complex ores.

Many deposits of complex sulfide ores are worked and at least some of thefmetal values extracted therefrom on a marginal Vcommercial economic basis. Some complex ores are so diliicult to handle that considering the recoveries and the cost thereof they are economically unworktot Patented Jan. 18, 1966 able under present conditions. In general, even in the case of relatively-simple complexes such as chalcopyrite,` there is a complete failure to obtain all of the metal values and frequently the costs are far out of line with- 'whatthey should be from an economic standpoint.

The principal object of my present invention is the provision of a new method plex ores. v

Another Objectis to treat a complex ore in such a manner as to break the complex so that the ore or concentrate may then be treated in accordance -with any one of many different chemical VandV metallurgical processes.

Another object is to recover all of the metal values of and means for treating comfrom complex ore including substantially all precious metals and rare metals -present in an over-all process which does not require selective mining or` selective concentration at any time. Y

, yAnother object is to recover all of the metal values from ya complex ore by a physical-chemical mechanism which permitsA production of a bulk concentrate of the vore as one step in its treatment in which bulk concentrate all of the metal values may be found.

Another object is the provision of highly concentrated l y specific metal concentrates from complex ores without leaving appreciable metal values in tailings.

Another object is the provision of means for the direct electrolytic recovery of metals such as'copper from ironcoppe'r sullides without the interpostion of a 'smelting procedure now commonly used in substantially all copper recovery methods.

A further object is to concomitantly secure substantially all ofthe metal values from complex ores and re-` to a relatively low temperature roast under controlled K combinations thereof for the recovery of substantially all of the metal values, including values of precious metals v atmosphere conditions. in such a manner as to break the complex comprisingthe ore and make it available for subsequent processing in accordance with substantially any known chemical and/or metallurgical procedure, or

and rare metals; i

FIG. 2 is an illustrative flow sheet showing one manner in whicha complex ore containing iron, copper, zinc and lead can be treated in accordance with the present invention for'V the -hydrometallurgical recovery of all of the metal values therefrom;

FIG. `3 is another illustrative flow sheet in which a 'l bulk concentrate of a complex ore is first roasted in a controlled atmosphere, the iron removed therefrom, and

the residue then subjected 'to flotation to grade concentrates oflead and zinc;

FIG. 4 shows one manner of treating a bulk concen trate of a complex ore comprising essentially iron copper sulfide wherein a copper sulfate solution is formed 'and suchsolution delivered to an relectrolytic cell for electro produce high recovery of the copper values, the tiow sheet also' showing other methods of treating bulk concentrates of complex ores comprising essentially iron coppersuldes, and

' `FIG.I 8 shows still another flow sheet for the hydro-` ton I first prepare a'bulk ,concentrate of the ore to `be treated, mining the ore as it comes without leaving unmined oresA containing diicultly removed impurities, such as is a common practice in the pyrornetallurgical treatment ofcomplex ores comprising essentially iron-copper in various pockets of the ore will be recovered, I may treat substantially any fraction of a ,complex orein accordance. with my invention and beneficially recover trough separatesthe furnaceinto two parts, a lower part 12 containing preliminary heating elements 13 and an 4sullidesv.v While preferably I utilize a bulk concentrate to be assured'that all of the metal values in theore and metal values which aretnot now normally recovered in i commercial practice. Illustratively, concentratesof ironcopper sulfide prepared for smelting will frequently leave someof the metal values in the mine or in vthe tailings, but I can treat such a concentrate in accordance with my present invention and obtain a better recovery of the metal values at less4 cost than is possible in accordance with conventional smelter operations. As a further illustration, I may, for example, treat a so-called ,selectively floated concentrate sold commercially as eithera zinc concentrate, a lead concentrate, or a copper concentrate, and recover metal values from the ore in greater percentages and at less cost than are usually' possible with present methods. When in the explanation hereinbelow, therefore, and inthe examples given, reference is made toa bulk concentrate, this is intended to` represent the most desirable procedure, but it should be borne in mind that concentrates representing something other all of the metalvalues-in the ore can also be treated inv accordance with any of the processes of the present invention." t' f L In accordance with the general featuresofthe inventionI iirstheat the ore vor a concentrate of the ore at a temperature between about 370 C. toaboutfOOf C., but preferably between about 400` C. and 550 C., in-an atmosphere comprising a minor,proportion of air and a major proportion A of-an. inert gas preferably including SO2, as will be explained, for a'period of from about three quarters ofan h ourto about one and one-half hours,

or until the iron content ofthe ore is soluble in ldilute sulfuric acid, the roasted product s o pr'mluced then being characterized by the fact' that the physical and/orchemical complex which characterizes vthe mineral comprising.

the ore is now broken and the metal values can as a rule then be separated -by any of several known methods including leaching, notation, and other 4ltnown procedures. kBefore referring more in detail to the parameters-'of the controlled atmosphere roast ,`the schematic illustration'of equipment and method shown: in FIG.I 1 will first be discussed. The data shown in FIG. 1 is based on-a pilot plantunit. through which approximately a ton `a day of ore may be processed, the exact amount processedv depending upon'severalfactors including the percentage of iron in the orebeing treated, the temperature, the furnace atmosphere, thespecic roasted end product' desired, etc.

In FIG. 1 I show a long furnace 10 which for all practicalpurposes is Vsealed except for passageways t o .be described for the admission of gasesthereto, theremoval of gases, theintroduction of the ore to be roasted (the term ore is used tov include either the ore itself or anykind of a concentrate thereof). A bed 11 which may comprise a strip of stainless steel in the generalform of a Shallow upper portion 14 containingthe ore4 and the furnace at,- rnospherein contact with the. ore. The ore is delivered from a bin 16 suitably sealed lto prevent entrance of air through a conveyor 17 which delivers the ore in a'continuous controlled quantity stream. In the drawings 17 isshown merel'yas a chute, it being understood that any one of several different types of feed mechanism such as a screw conveyor-may -be employed. n

The ore is delivered to the bed 11 and is then continuously and slowly advanced alongthe bed by a screw conveyor 18. Preferably this is a series of overlapping paddles rather than acontinuous screw, so' that as the ore is advanced it is also turned over and fresh surfaces continuously presented into contact with the atmosphere. The roasted oreis discharged at the right hand side of thev furnace, looking at FIG. l, through a'cooling and sealing chamber 19 and thence to asealedrreceiving bin 21, the roasted ore thereby being kept out of contact with atmospheric'air'while' it is in a heated condition'.

`Oxidizing air is delivered to the furnace4 at at least one location between its ends, as indicated by the -two air inlets 22 each of whichv is controlled by a valve '23. The number, location, disposition and controlled size of the ports foradmitting the air are determined by a number of factors, as vwill be apparent,including the amount of inert gas introduced, the proportion of inert gas to air des'iredand required, the proportion of metal such as iron tobe solubilized at a given treatment step, and other v factors.

The hot'gases from the furnace are discharged-through a stack-24 which .runs to any suitable location as determined by an over-all design,-.but in Aaccordance with the present invention optionally toa sulfuricacid plantl for theproduction of sulfuric acid usable in subsequent steps of the process. The stack gases from equipment such as shown` in FIG. 1 are particularly suitable for the producn tion of sulfuric acid because all of the arsenic in the ore is removed with the iron and practically none of 'it is 'driven off during the first heating step. As aconsequence,

there is no'a'rsenic to poison a catalystas frequently occurs in sulfuric acid plants where the sulfur dioxide re-v sults from high temperature treatment of ore bodies or fractionsther'eof. y

A blower 26fwitlidraws stack gases from thejstack 24 through a pipe 27, the stack gases first going through a dust ccllector`28, thence through a'heat exchanger 29 where the gas is cooled to a pre-'determined temperature usually about room temperature or somewhat below, and thence delivered to or through `a reservoir 31 and. through an inlet pipe 32 in controlled amount `to the discharge vend ofthe furnace. The vflow of cooled and cleaned products of combustion tothe reaction zone may be con'- trolled in various ways,` and to indicate the controlI show a conventional valve 33. For starting the'operation sulfutr'dioxide may be produced in a generator 34and delivered through a pipe 36 and valve37 v to the container 31.

The temperatures and broadly indicated' atmospheric conditions are ,typical of the first step inthe treatment of acornplex ore containing substantial amounts of iron sulde whenthe purpose of the first step isto solubilize allof the iron and break .the sulfide complex. 'The conditions indicated are representative and mayvvary, particularly if Athe FIG. 1 equipment is used to solubilize other metals, as discussed later on inthe present specication; Itv-will be' noted that at each-.end of the furnace the furnace atmosphere consists essentially of only nitrogen and sulfur dioxide and that in the center portion of the furnace there lis a mixtureof nitrogen, sulfur dioxide and oxygen, it being obvious that in `general the oxygen content is apt tov be somewhat higher in the immediate area where the air is introduced than' further to the left thereof when there has already beenfsome reactionbetween the oxygen and ore., The temperatures given are for the ore rather than for the hot gases above the ore. The ore enters generally at about room temperature and gradually increases as it moves toward t-he discharge 'end of the furnace. y

It will be noted that a temperature as high as 550 C.

may be present at the extreme end of the furnace toward v the discharge end, and that normally a temperature of `impurity4 in` chalcopyrite and a concentrate produced this magnitude will tend to produce some insoluble fery in. the stack gases clearly indicates that the mechanisnt involved is not one in which ferrie iron is first produced and the ferrie iron then reduced by means of sulfur dioxide. There is something in the physical chemical equilibrium involved of which l am not fully aware which causes the production of ferrous oxide to be complete, and widens the temperature range at which ferrous oxide can be produced, particularly in the direction of raising the temperature above that normally possible in this type of operation and as disclosed in my prior patent and-copending application. v Y.

Before describing the various examples of the invention as shown in the ow sheets in detail, it may be helpful to consider the processing of the present invention from a general standpoint.

I have already referred to the fact that in a single step all of the iron in the roasted product is in the ferrous condition,l all of it apparently being present as ferrous oxide (FeO). The ferrous iron may be leached and recovered in the form of a salable product in accordance with several diierent procedures. Arsenic present in the containing the selenium, a selenium-copper alloy will be formed during `the smel'ting procedure and thereafter it will Lbe a physical impossibility to extract the selenium from such copper. Such GOPperis then for all practical purposes non-usable, and there is more than one smelter in this country which has a pile of vpigs of blister copper containing selenium which in the present stage of technology cannot be puried and which represents a. dead loss.

Another feature of the present invention is that I may take a complex ore, treat vit to break" the complex,'and then by selectivev flotation produce'very much higher grade concentrates than can be produced by previously known methods, and the concentrate so produced may be sold as such to commercial companies at a `good price because of the mineral content and without penalty of the type charged when the concentrate includes more than a xed amountv of undesired elements. may then be further treated by, for example, conventional smelting-procedures, by means of electrodeposition, or any of the several usual procedures common to chemical and metallurgical operations; When the concentrates are treated by smelting operations, there will be some loss-of precious metals and usually a complete loss of rare metals subjected to the sinelting operation with the concentrate. In the preliminary roasting step, however, prior to Athe production of the concentrates, I may recover certain of the rare metals. of the selenium present vaporizes and passes to the stack with the eiiluent gases, and the selenium so'vaporized is readily recovered by simple precipitation methods.

In the specication and particularly in the claims I refer to the treatment of complex ores or complex sulfide ore is found in the resulting ferrous sulfate solution ap- 1 parently a's an iron arsenite, and it may be completely removed by adding metallic iron in the form of very fine particles, the arsenic then being precipitated as elemental arsenic.

A very important feature is that gallium which is present in fair amounts in many ores, but generally not recovered at all, is leached out with the iron and is completely recoverable. y

Most sgnilicantly, the rst step in accordance with FIG. l breaks the complex so that the various metal values can be separated in accordance with many different schemes including flotation. While theiron is preferably removed by leaching, it also can be concentrated and a fine separation produced by flotation means.

Still another feature o the method, particularly including the controlled atmosphere roast, is that precious metals and particularly rare metals can be recovered with substantially no loss. They are recovered in various ways as will be pointed out, but at this time it may merely be tical purposes lost. Illustratively, when rare metals are present in small amounts in concentrates fed to copper smelters, the rare metals are either slagged otfwith the iron silicate Vslag or with the slag from the reverbcratory furnace, or they may be driven olf with eiuent gases and lost to the atmosphere. More than a certain amount of certain rare metals cannot be tolerated. For example, if selenium is present in fairly substantial amounts as an ores, and it is in'poin't to explain at this time that by a complex ore l mean one in which the metal constituents are complexed in the mineral with sulfide radicals in such a way that the metals cannot be separated by usual means, particularly means involving physical manipulation such as tabling, flotation, sink-float, and other such mechanisms. These ores are. in general 'sulfide ores, but they may have a small inclusion of oxides, carbonates and the like resulting from partial weathering of the sulfide.

When I employ the term ore it is in a generic sense` unless' the context indicates otherwise, and includes any4 amount and complexity ofthe equipment required to recover metal values from complex ores is greatly reduced. I may, for example, deliver an ore or concentrate to the reaction zone in the step ofv the process indicated in FIG. l with particle sizes up to one-quarter of an inch, although to produceeven a bulk concentrate it is usually necessary to comminute the' ore `to a point below this' size. In any case, whatever the size `of the concentrate delivered to the reaction'zone ofthe controlled atmosphere step, the product is decrepitated and comes out as a very-fine powderlike mass. In general, therefore, when operating in accordance with the present invention, crushing and grinding equipment can be reduced by at least about fty percent and up to seventy-tive percent in uuus'ual cases, and the number of otation cells employed can also at least be cut in half. It should be apparent that when a selective concentrate is produced, the ore must as a rule be ground to about 300 mesh, and there must beseveral passes through different types of flotation cells before the end product concentrate is produced clean enough for commercial exploitation. This should The concentrate so sold For example, a substantial proportion t 7 :contrasted with the present invention in which it will try seldom be necessary to grind the ore to less than mesh, and in which the bulk concentrate may be 'oduced with a single battery of Aflotation cells with no :ed for loops to specially process certain fractions and fa rule with no need for the use of substantial amounts expensive frothing and collectingy agents. `In accord- [ce with the present invention such flotation as may be :orn'plished after the initial controlledtemperature roast a bulk concentrate requires no additional'grindingbeuse of the decrepitation in the roasting step, and a-s a .le very simple and inexpensive flotation reagents can en be used to produce the desired concentrates.

I have referred to the use of an inert gas or neutral In determining the cooled stackgases, for example, the temperature would gradually rise to a point where undesirable insoluble and v insolubilizing compounds were produced. One approach in 4solving'the rising temperature problem wouldbe to 'n greatly restrict the amount of lair admitted over a given period of time and, of course, it would also be necessaryl to greatly restrict the speed of movement of the ore through the reaction zone. Such a method might require twenty-fouror forty-eight hours or still more lto produce the necessaryflow temperature oxidation, as contrasted with, for example, one 'hour under circumstances illustrated in- FIG. '1.

v The second function of the added inert gas is the dilutionof the oxygen in the `furnace atmosphere,and the reduction ofthe actual amount of oxygen the atmospher'etlowing across the top of the ore. This results in part in=reducing the rate of oxidation, vbut it should be noted that the rate of oxidation is more properly a function of ambient heat rather than oxygen dilution, keeping' in rnind as previously pointed out that. there must be enough oxygen present over a period of time to accomplivsh the degree of oxidation desired, and this 'amount of ry-amounts, but must be calculated in accordance with ert gas is then established based on the predetermined rygen and airl requirements. The'inert gas introduced ay be of many types but preferably includes a propor- Jn of sulfur dioxide, all ofiy which will be 'considered uther in a discussion of the functions 'performed by the luting inert gas introduced. y

The'amountof inertv gas introduced may 'vary extenlrely depending upon a number oflfactors including the laracfter of the ore itself, and particularly the iron connt of the ore yv vh'en the function of the first step is to lubilize all offthe iron present, the temperature in the action 2onethe timeof reaction as measured bythe eed at whichl the ore passes through the roasting furlce, the type of inert gas, particularly the amount-of `lfur dioxide included, and other factors. vwever, there should be .a major proportion of inert gas ld a'minor proportion of air ranging all the way from out one part (plus) of inert gas to one part air, and up and even in excess of tenparts Iof, inert gasto one part The amount of air required In general,

' air. In determining the furnace atmosphere it should,

i course, be noted that some of the sulfur` the ore is Inverted, to'SOz-so that there is an SO2 fraction in the rnace atmosphere resulting from the roasting operation.

lien the .relationshipof inert gas to air is relatively the me, then the sulfur dioxide generated'by the roas'tbemes a greater factor and this'rnay be Vfurther accented l control of thefurnace atmosphere to retain e'iuent .ses fora longer than usual time before discharge rough the stack. As pointed out the introduction of inert gases into the rnace, and particularly the introduction of cooled stack tses into the furnacev in accordance with the preferred abodiment of the present invention, performs several nctions.

The' first and one of the very important functions of e added inert gas at ro'om temperature is a scavenging )eration `in a sense, but has to d o not 'only -with the reoval of the heatof reaction so that the temperature of e ore may be held down to a desired point such, for ample, as between 400 vC an'd 550 C. in thelhighest mperature part of the reaction zone. If the mechanism town in FIG. l'were operated without the addition of oxidation would still occur rapidly regardlessv of dilution if the furnacetemperature were allowed=to rise. So far as the dilution of the oxygen is concerned, the inert gas i might, for example, be free' of SO2 and could, for? example, be nitrogen. Excess nitrogen might 'also be introduced, of course, vby partial combustion of the oxygenin the air delivered tothe reaction zone, but in general this particular approach to dilution of the oxygen is not recommendedbecause of the economics involved and for several reasons'.` L

The third function of the introduction of an inertjgas isconce'rn'edprimarily with vthe sulfur dioxide content of the furnaceatmosphere, and in consideration of this function it is apparent lthat an ordinary inert gas without a sulfur dioxide content would not produce this particular function.v I have observed that while the sulfur dioxide apparently does -not enter the reactionvin any way as a reducing agent, its concentration has an apparent direct effect in preventing the formation of ferric iron which normally cannot be reduced at a temperaturebelow about 600 C., in breaking the complex, in permitting increased temperatures, and widening the temperature' range of a desired reaction, and` solubilizing one constituent of the ore such as the' iron without solubilizing other constituents. All of Ithe above, of course, is subject to proper control, but the results are apparently not obtainable exa fraction of sulfur dioxide almost act's like a catalystin Y.

making the controlled Y atmosphere roasting step more selective in that it permits confninginitial decomposition of the sulfide to a selected metal such a's iron. `This function was touched ori hereinabove, but bythe presence of the sulfur dioxidethe decomposition of the iron sulfideand, of course, the-concomitant'breaking of the sulfide complex-may carried out .at a higher than normal temperature (suchas disclosed in my prior patent and prior pending application) and overa wider range than heretofore possible while still confining the solubilizing to the single. metal or, in some cases, a group of metals as pre-determined. v

Another direct contributing function of the sulfur dioxide'in which it does not act as a reducing agent nor apparently technically as a catalyst, but with some apparent function like a catalyst, is the complete breaking of the complex comprising the mixture of metallic sulfides. As I'have pointed out, wheniron is converted to FeO and the sulfur equivalent in the iron removed, the

, ent invention,

remaining base metals appear then to be in the form of simple suldes exemplified by CUS, CuzS, PbS, ZnS, CdS,

etc., all of which are then readily separable by severalv sion of the iron sulfide to FeO, although so far as I cany now determine substantially concomitant therewith. I

have reason to believe from certain observations that with suitable controls it may be possible to break the complex without converting the iron, or at least all of it, to Fe() if there shouldbe any reason for doing so, and, therefore, I do not limit myself to specific conditions for iron removal as exemplified in the present specification.

Another apparent function of the controlled atmosphere roast, which may or may not be directly associated with the SO2 content of the furnace atmosphere, is that the lirst roast and iron leach appear to pre-condition or activate the residue remaining after the iron leach for a later high temperature or Vthe like roast in that the normal high temperature roast can then be carried out at a lower temperature than normally required and/or for a shorter period of time. v

Still another function or feature of the controlledat` mosphere roast is the ability to separately solubilize varions constituents ofthe complex ore for separate leaching operations if this manner of treatment should be required for any particular reason. For example, after the iron has been removed I may then take the residue from the iron removal end and again pass it through the reaction zone, as shown in FIG. l, keeping the conditions discussed standard except for raising the temperature some,4

illustratively to 550 C. The residue can then be treated to remove the lead hydrometallurgically, as, for example, by first treating the residue with sodium carbonate to form lead carbonate, and then leaching the lead carbonate with acetic acid. ln any event, after the removal of the lead the residue can then again be run through the controlled atmosphere roasting procedure at a temperature somewhat higher,l illustratively about 600 C., and the copper selectively solubilized. After the leaching of the copper, the residue can again be treated by the controlled atmosphere roast at, for example, above about 650 C. and the zinc then removed by leaching. Illustratively also as will be pointed out, two or more of the metal constituents can frequently be removed in a single roasting leach step if required, either in the first step of the process or in a subsequent step. be roasted to solubilize only the iron or, under certain circumstances at least, the roasting conditions can be controlled to solubilize both the iron and copper.

With further respect to the percentage 'of Vcooled stack gases re-circulated back tothe reaction zone, I have already pointed to some of the variables that have a bearing on this percentage, and I wish to note again that the physical characteristics of the ore being treated kand the amount of stirring in the reaction chamber also have a bearing on this percentage, in addition to the previously discussed variables including the temperatures, the time of reaction, and the chemical characteristics of the ore. While some of the examples will illustrate the air/inert gas relationship within the general range of about one to one to one to ten as already discussed, I wish to point out that a common percentage for many types of complex ore is of the order of four, five or six parts 'of cooled stack gases by volume to each air volume. As an example, a complex: ore from northern Arizona containing about twenty percent'iron, twenty-tive percent zinc, about ten percent' lead and about two percent copper required about six parts of cooled stack gases to one part of air for best operating results in a relatively small pilot plant in which the ore Illustratively, chalcopyrite may .that the iron is recoverable as asalable product.

ld passed through lthe reactionzone in about one hour or slightly more, and inwhich the total amount of ore treated in the reaction zone for a twenty-four hour period was slightly lover one ton. On the other hand, a high zinc ore containing also a fair percentage of both lead and copper but only a trace of iron required only about one part of stack gases to one partof air to solubilize all of the iron and leave the zinc present in the form of a simple sulfide. I have alsotreated a concentrate comprising essentially iron disulfide to determine the reaction conditions for solu- -bilizing all of the iron therein and, with substantially the `treatment being at about onequarter of a ton per day.

' One of the features of myV invention is that with a'single roasting step l may solubilize all of the iron present in a complex ore and remove this iron in soluble condition in accordancewith several differentl procedures. From the standpoint of iron recovery it must be remembered that in almost all ore treating operations iron is treated as Van impurity which must be removed, and it is very seldom l When concentrates of lead, rrincer the like are sold to processors a penalty is charged for an iron content over an established ligure to thus reduce-the net price which the commercial smelter or other processor pays to the producer of the concentrate. As a result, producers of commercial concentrates normally avoid as much iron as possible in their finished products, even at the expense of loss of otherwise valuablev metal. The iron removed goes to tailings and, generally speaking, has no commercial value. rConcentrates of iron-copper sulfide provided 'as a feed to copper smelters will run as high as forty parts of iron to twenty parts of copper, and in some instances even the copper in .the concentrate is of a still lower value. This necessitates repeatedly charging converters with new charges of matte before it is possible to build up suflicient copper sulfide in the .converter to warrant continued operation' and Bessemerizng of this material to the usual blister copper. As an example, it is frequently necessary to provide as much as twenty or more charges, taps or ladles of copper matte for each nished charge of white metal with, of course, attendant losses of time and increased labor costs.' But most significantly, all of the iron in the' concentrate must be handled several times in its path from the minethrough the smelter, and it ends up on a slag pile in the form' of a product consisting essentially of iron silicate and for all practical purposes is completely worthsalvagin'g this tremendous mass of iron, but none of this development work hasV resulted in even 'a vestige of success.

In the practice of my invention I recover the iron almost as a by-product and almost without any cost directly attributable to iron recovery. Normally the iron is leached as a ferrous sulfate solution, and as a rule represents a pure iron sulfate except for the readily removable arsenic and some small amount of antimony in the particular ore treated. Later the ferrous sulfate may be crystallized out of solution, o'r the solution dried on commercial drying equipment to recover a commercially valuable hydratedferrous sulfate' having such extensive de'- mand in the agricultural fields that tremendous quantities thereof are imported into thiscountry. Another method of recovering the iron is to heat the ferrous sulfate in a non-oxidizing atmosphere to drive olf sulfur as SO2 and with some S03 to form a product' consisting essentially of FeO. The FeO can be the starting point for producing many different types of commercial iron compounds to metallic iron at red heat by means of dry H2. The

lll

finely divided metal is very valuable for many purposes .ncluding, for example, precipitation of arsenic from the ferrous sulfate solution, the precipitation of antimon'y and for the removal of impurities from solutions by means of 1n oxidation reduction process in which the iron goes into zolution and the reduced metal precipitates out. The inely divided iron so produced is very valuable as a -re' Jlacementy for old cans andthe like frequently used in vhe copper industry for replacement of copper in solution ts in vlaunders and the like. Still another method which nayv be used withv the ferrous sulfate is to heat it in air tt about 600 C. to drive olf S03' which may then be iydrolyzed to form sulfuric acid and to produce from the 'errous sulfate a ferrie oxide (Fe203) which also has nany uses in agriculture, in the pigment industry, and the ike.

Wlhile various complex ores' have definite similarities :ven though taken from different locations, it is very eldom thattwo -ore deposits will be absolutely identical. legardless of detailed instructions which may be given, herefore, with respect to treatment of a given ore, it naybe Ithat a closely similar ore will require experimenation and even quite changed conditions for most effeci`ve use `of my present invention. The present specificaion including the detailed examples, however, will furnish tn ample guide to those skilled in the art for changing he process as required. There may be, however, other :o-nsiderations not directly related to the present invenion, but having' a bearing on most effective use of the resent invention. It is impossible to give a complete iibliography with respect to this subject matter, and I vill attempt only one illustrative situation. Complex ulfide ores frequently have a content of alkali metals ralkaline earth metals, and the presence of either, paricularly in relative-ly large amounts, may have a delfteri'ous effect on the present process. By firstv washing ,nd leaching the ore .or preferably its concentrate by neans of dilute sulfuric acid, base constituents such as odium, potassium, magnesium and the like are readily e-moved. The first roasting step of the present invention then more effective. I do not mean to say that without he preliminary treatments with dilute sulfuric acid that t is impossible to operate the present process, but as .rule it will be found that the time of treatment Will have o be increased somewhat and the temperature also'in-l reased somewhahif the iron is to be completely soluilized, and,v moreover, `iron solution will then be coni'minated by at least some of the alkali'or alkaline earth ietials present.

Concerning the. rte-circulation of sulfur dioxide or the irect introduction of sulfur dioxide from a sulfur dioxide enerator, or such other manipulation as 4may be emloyed to increase substantially tlhe sulfur dioxide in 1e atmosphere of vthe'reactio'n Zone. I am aware 'that hysical chemists studying the reaction occurring in coperconverters Where the temperatures'w'ill range from 0.00, F. Ato 2600` F. A'have attempted to explain the actions occurring in the `molten copper and tihe equil brium involved by reference tothe partial pressure of Jlfur dioxide above the molten copper or -copper sulfide. lass. Therehave a1so been some references, althoughot so precise, with respect to partial pressure of sulfur ioxide in studies relating to what are frequently called Iaight sulfate roasting techniques. These techniques lvolve roasting a material which may, for example, be sulfide' ore at relatively Ihigh temperatures, such as C. to as high as 110() C., for the purpose of con'- erting some of the constituents yto oxides or 'sulfates 'hich can then be leached from the roasted material. 1 accordance with known sulfate roastingl procedures in-v olving high temperatures it is, of course, possible to l)lubilize some of .the materials present, but as a rule as jindicated at 42.

million.

, obtainable by my present invention.

. 12 with respect to partial pressures of SO2, it appears obvious to me that the mechanism involved-and the explanations thereof incopper smelting procedures and sulfate roasting procedures donot account for the unusual results i While it may be that the partial pressure o f SO2 may be-found to have a vbearing on the reactions -occurringbetween the ore on the one `hand and the furnace'gase's on the other and, of course, Iwithin the ore itself Where' there is an obvious re-arrange'ment of chemical bonds and possible loss ofl sulfur other than that accounted for by losses from the ironv portion of the complex, explanations made heretofore with respect to other reactions do not account .for

'results in the present inventionQ I am unable to explain the results which I obtain from a technical physical chemical standpoint, except that it appears from the rela-` tively small proportions of S03 in the stack gases that the action of the SO2 is not explained -merely by assuming that it functions as a reducing agent.

l I Example I In FIG. 2 I illustrate one general method Yof practising the present invention assuming' the original ore to be a complex sulfide ore containing substantial proportions of iron, copper, zinc and lead lwith or without minor lproportions of vother materials such asselenium, gallium, arsenic, antimony, cadmium, indium, bismuth and the like. In accordancewit'h the'exampleof FIG. 2 I produce a bulk concentrate of the ore atl 41 and subject the ore to a low temperature controlled atmosphere roast The vc onditions ofthe roast` are indicated generally at FIG. 1, with four to six parts of stack gas cleaned, cooled and delivered back to the ore delivery end of the roasterv to one lpart of outside air, the exact amount of re-cir'c'ulated and cooled 'stack gases being determined by the proportion of iron in the ore with about six`- parts of stack gases vbeing r'e-circulated when the iron content ofthe ore is of the order-of fifteen to-twenty percent. The time of movement of a charge through the reaction zone is about one hour. Between 'seventy-tive and ninety percent -of the selenium present in. the ore is recovered from the stack 'gasesat 43. As

recovered, tihe selenium exhibits a reddish color but boil-l ing in clear water will convert it to the usual selenium grey. Its purity is very high, the total proportion of impurities being of the order of ten to fifty parts per The roasted ore is received from the reaction zone at 44 and is allowed to' cool to room temperature before being brought into contact with the air, On examinatio'n it will be found that the complex is broken andthe metal constituents present are al1 in the form of separate compounds, 4usually a mixture of iron oxide and sulfide of.- -the non-ferrous metals, with apparently some traces of sulfate The roasted ore is then brought to a leach station where it -is leached 'with vdilutesulfu-ric': acid and ayleach solution recovered at ecoveries are low and may even be less than sixty percent 47 containing ferrous sulfate with all of the arsenic in the ore in solution probably as an acid, with a substantially large proportion-of the gall-ium and some of the antimony,` The leachfsolu'tion is then treated at 48 for the removal of gallium, arsenic and antimony, and the production of a substantial-ly pure ferrous sulfate solution at 49. A suitable purification procedure is to treat -the solution .at 48 withstoichiometric proportions of nely divided ir-on powder to precipitate out arsenicanrl antimony, both as the metal. Galliurn may be precipitated at tihis point as shown at station 51, or the gallium may go with the ferrous sulfate solution and subsequently be removed from such solution in accordance with several different available procedures.

. At 52 the' residue remaining after the acid leach to recover iron is subjected to a relatively high temperature oxidizing roast above 570.C. and suitably as'high as 650 C. to 700'* C. to solubilize remaining constituents by driving off remaining quantities of SO2 and leaving the materials in the form of oxides or sulfates, depending upon the equilibria involved. When the high temperature roast is employed, the balance of the selenium in the ore will be found in the stack gases and can be recovered by suitable precipitating means as indicated at 53. After the high temperature roast, the residue is subjected to an acid leach at 54 to remove zinc, copper, cadmium, indium and the remaining gallium to leave a residue at S6 which contains all of the lead, the gold and silver and such bismuth and germanium as may be present. Antimony if present may be found either in the residue at S6 or in the sulfate solution as at 57, but commonly will be distributed bctween the residue and leaching solution. The leaching solution contains sulfates of zinc, copper and cadmium and indium if present in the ore. These values may be recovered in any suitable manner as indicated at 58 in cluding any known chemical and hydrometallurgical method, or in accordance with any one of the procedure discussed in my issued patent or co-pending application. lllustrativeof basic techniques are the precipitation of copper with metallic zinc to increase the zinc sulfate content of the solution and precipitate copper out as finely divided metal particles. Cadmium and indium may both be removed from the solution by adding additional amounts of tinely divided metallic zinc, the cadmium and indium being precipitated as metals. In each instance there will be some metallic zinc mixed with the metallic cadmium and indium, and the mixture can be set aside for later treatment to produce an effective separation. It is possible to separately precipitate the cadmium, but sometimes there is nothing gained thereby. Zinc can be stripped from the zinc sulfate solution electrolytically or the zinc may be recovered by any one of several known chemical procedures.

The residue at 56 from the acid leach of the zinc, copper, cadmium, etc., from the high temperature oxidizing roast contains all of the lead, such gold and silver as may be present, and also such bismuth and germanium as may be present in recoverable condition. The lead is lirst removed by converting it to lead carbonate as indicated at S9, and then leached out with acetic acid to form a lead acetate solution as shown at 61. The lead acetate solution may be treated with sulfuric acid to precipitate the lead sulfate and the acetic acid solution regenerated for further leaching of lead carbonate. To form lead carbonate the residue is normally treated with a slurry of sodium carbonate, and the resulting sodium sulfate formed by conversion of the lead sulfate to lead carbonate is withdrawn as a sodium sulfate solution as indicated at 62.

.The residue as shown at 63 is now a highly concentrated solid mass, as contrasted with the original ore, so that even though the gold and silver and bismuth and germanium if present were only present in trace amounts in the origf inal ore, they will be found in fairly concentrated amounts in the residue at 63. The residue is treated with a cyanide solution at 64 and gold and silver cyanide recovered at 66. The residue containing bismuth and germanium is delivered to 67 and leached with a twenty-five percent hot sulfuric acid solution at 68 to produce a solution at 69 comprising bismuth and germanium sulfate with an excess of sulfuric acid. This solution is then suitably treated to recover bismuth and germanium. The acid leach may leave some germanium in the ore which may be removed by a sodium hydroxide leach, or an alkali metal hydroxide with some excess of the hydroxide used directly to leach the germanium. The tailings at 71 comprise essentially only SiO: and the like gangue materials. By weight it will normally represent approximately 0.15% to of the original heads, that is to say the starting bulk con- Gold, oz. 0.446 Silver, oz 13.6 Iron, percent 27.4 Zinc, percent 28.1

Y Lead, percent 12.9 Copper, percent 1.05 Cadmium, percent 1 0.11 Antimony, percent 1.02 *I Arsenic, percent 1.3 Bismuth, percent 0.007

have not shown, for example, the common practice of returning stripped solutions for further leaching, the buildv ing upof residues of trace elements and the recovery of the trace elements after they have built up to a reasonable value, and many other details which are characteris y Example 2 Proceeding generally accordance with Example 1 and the flow' sheet of FIG. 2, a bulk concentrate was made of an Arizona ore having the following assay:

Gold, oz. 0.375 Silver, oz 11.11 Iron, percent' 22.78 Zinc, percent 24.29 Lead', percent 10.6 Copper,'percent 0.865 Cadmium, percent 0.09 Antimony, percent 0.87 Arsenic, percent 1.2 Bismuth, percent 0.006

The above bulk concentrate contained:

Pounds of iron 455 Poundsoffzinc 485 Pounds of lead 212 Pounds of copper 17 Pounds of cadmium 1.8 Pounds of antimony 17 y Pounds of arsenic 24 Pounds of bismuth '0.12

The listed materials totalled slightly over 72% of the original total of 2,000 pounds, and the balance comprised sulfur, silica, some alkali and alkaline earth metals, and some rare vmetals including gallium, indium, germanium and selenium. The'assay given was a commercial assay and did not identify .the presence of any rarel metals, but in the final recovery I found substantial percentages of such rare metals. l

`The concentratewas then roasted without preliminary removal of the alkali and alkaline earth metals for two hours at a maximum temperature of 450 C. and with a furnace atmosphere characterized by the introduction of one part ofair to six parts of cooled stack gases. The roasted concentrate was then reduced to 1,920 pounds and it assayed as follows: f

The elements `listed above comprise 85.9% ofu the total,`

furic acid and the solution from this leach was found to contain:

Pounds of iron as the sulfate 470 Pounds of arsenic 26 Pounds of antimony 2.9

The solution was agitated with elemental iron powder and the arsenic and antimony precipitated as metal.

. nesium as sulfates.

^ l After precipitation of the arsenic and antimony with iron the remaining solution was found to contain 506 pounds of iron and a small amountof calcium and mag'- This solution was then put through an evaporator and crystallizer and the end .product ofv the operation consisted of 2450 pounds of hydrated ferberemoved, and'this may be important depending upon specific control desired with respect to the residue after iron removal. y

'Ihe washed and dried residue after the acid'leach to remove the iron weighed 1060 pounds. It assayed as follows:

Gold, oz. 0.693 Silver, oz. 20.6 Iron, percent 0.4 Zinc, percent --..V 46.0

In theabove assay vthe figures for gold and silver are given Vin terms ofounces per ton in accordance with usual practices, but if these are calculated in terms of ap proximately half a tonthe figures-would be -.'.Mounce Lead, percent -s u 21.0 Copper, percent 1.6 Cadmium, percent 0.167l Antimony, percent 1.37 Arsenic, percent Q Nil Bismuth, percent e 0.011

The copper is separated'and'is found to be very pure and can be employed directly in .producing many types of copper hardware or, of course, alloyed with zinc, vtin or the-like to produce brasses and bronzes of commerce.

After the removal of copper in the manner described, thesolution was further treated by the addition of five pounds ofzinc powder and agitated until the solution was substantially neutral, that is to say had a pH of about 7. Under these conditions the Y'solution will contain some freezinc which, of course, will be `precipitated in due course with the cadmium. The precipitate was found to contain sixty percent cadmium and forty percent zinc. In actual practice this Aprecipitate is preferably put to one sidekuntil asubstantial quantity `is at hand, after which thelzinc and cadmium can be separated in accordance with one of several procedures, none of which is particularly concerned with the, present invention. Cadmium represents a'metal in relatively short supply, and, a1-

though it is not technically classified as a rare metal,

t several chemical and/or hydrometallurgical methods for recovering zinc values from this solution none of which is particularly peculiar to my present invention.

per half ton of gold and 10.3 ounces per half ton of silver,

which, of course, is generally in line. with the original assay figures. p l This residuewas roasted half hours. was carried on for one hour at a temperature. between 700 C. and .750 C. The results were about the same in each instance,and the further explanation which follows canassume that the roast occurred -in connection with either set of conditions. The roasted product in each instance weighed 980 pounds. The assay was substantially the same as given immediately hereinabove, except that the percentageswere somewhat higher because of the'replacement of much of the sulfur driven olf with oxygen. In this particular case practicallyl all of the metals were in the form of the oxides with some slight amounts 'of sulfate. The conditions can be', conat 635 C. for one and one 'titolled so that the sulfatos predominate or represent substantially the entire mass, and, of course, itis understood that in general thel sulfates are as readily leached as the oxides. The 980 pounds of roasted-residue was then leached with dilute-'sulfuric acid ending with a pH between 2 and 3, or, in other words, with a very small amount of free sulfuricv acid remaining after the leach. 'Ihe leach solution contains substantially al1 of the zinc, cadmium. and eopper'with a trace'of iron.` The amountA of ironwas so slight, particularly when the iron leach iscouducted with some .grinding tobreak up `any cocooned particles, that it is very dicult to detect by any ordinary laboratory methods. An analysis of this l'each solution shows it to contain 480 pounds of zinc, V17 pounds of copper and 1.8y

In a modiicationof the process roastingv 'l The residue remainingafter the acid leach for the re moval of zinc, copper andl other metals was then washed and leached with twenty percent sulfuric acid at approxirnately 200 F. for-one hour, andthe leach solutionthen separated from` the solid residue. This leach solution contained 12.8 pounds of antimonyas the sulfate. I referred previously to a grinding leach to remove iron and in this connection it maybe noted that if such grinding vleach is employed, not only will at least- 99.99%' of the iron be removed, but a substantial amount of the antimony and sometimes all of -the antimony may be removed by this rst leaching step. It is-the antimony which is no tremoved during the first leaching step which visl removed by the hot concentrated sulfuric acidv leach referred to immediately hereinabove. By adding eight pounds of zinc metal lpowderl to the antimony sulfate Y solution a precipitate is obtained comprising antimony metal. From a solution containing 12.8 pounds of antimony Iv readily removed 12.1 lpounds of antimony metal by .the precipitation procedure discussed, and this perlone hundred and ten pounds of sodium carbonate yand the mixture heated for half'an hour fat 200 F. This treatment convertsthe lead sulfate in the residue to lead carbonate and forms sodium sulfate which is withdrawn vfrom the residue in the form of a solutionand the dry residue washed to remove all of the sodium sulfate. The residue is then .leached with a ten to twenty percent solution of acetic acid for approximately one hour at ordinary room temperature to convert the lead carbonate into lead acetate. The resulting lead acetate solution is withdrawn and one 'hundred pounds of concentrated sulfuric acid added toit to precipitate insoluble lead sulfate. The washed and dried lead sulfate weighed three hundred and ten pounds and, of course, the regenerated acetic acid is returned to storage for further use in leaching lead carbonate.

'the residue after the removal of the lead weighed always happens, and it appears to indicate that in some,

manner or other traces of gold and silver are so locked up in the complex that they are not detected by ordinary assay methods.

After removal of the gold and silver the dried residue was found to weigh forty-live pounds. It assayed:

Percent Silica 74 Iron 15 Antimouy 0.09 Bismuth s 2.8 Lead Trace and no zinc or copper. If desired, of course, the bismuth may be separated by suitable means from this residue.

Example 3 A bulk concentrate of a complex sulfide ore from Arizona assayed as follows:

i8 Example 4.

bulk concentrate was made of la complex Arizona ore taken from the Bradshaw area, A complete assay of the concentrate was not made, but it was determined that it contained:y v

About 1.12 oz. gold per yton About 33.3 oz. silver per ton About 30% lead About 2V2% copper About.35% zinc In excess of 20% iron Looking at FIG. 3, the bulk concentrateindicated at 76 was subjected to a controlled temperature roast as indicated. at 77, the ratio of stack gases to air being five to one, the maximum temperature 500 C., and the total time of exposure of the concentrate to the heat and atmosphere of the reaction zone seventy minutes. The initially i roasted ore after cooling was removed at 7S and it was found on inspection and testing that the complex iron.

zinc-lead-copper complex had been broken, that the iron was substantially all in the form of FeO, and that the` principal remaining .metal values, particularly zinc, copper and lead,`v were apparently in the Aforrn of simple sulfides exemplified by the formulae ZnS, CuS and PbS. The roasted-concentrate wasthen leached at 479 with a dilute sulfuric acid and a solution obtained at 81 comprising Au, oz. per ton .4 Ag, oz. 17.7 Zn, percent 35.5. Pb, percent 14.4 Fe, percent 10.5 Cu, percent 0.94 Sb, percent 1.3 Bl, .09 Cd, percent 0.11 As, percent 1.7

A ton of this concentrate was treated in accordance -with the process shown in FIG. 1, the time of treatment being one hour, the maximum temperature 470 C., and the feed back of cooled stack gases six times in volume that of the air introduced at the sides. yAfter the first roast the product weighed 1822 pounds, and after the first leach the residue weighed 1484 pounds. The Aresidue from-the iron leach was dried and heated in an oxidizing atmosphere for one hour at a temperature of 700 C. After such roast the Aproduct weighed 1394 pounds. After the second roast the mixture was leached with sulfuric acid and after the leaching the residue weighed 460 pounds. The zinc so recovered as zinc oxide and weighed 824 pounds, the copper was recovered from the solution as metallic copper and weighed 18.66 pounds.` 2.12 pounds of cadmium were recovered. The residue'after the second leach was treated in accordance with the process described to remove lead, and 420 pounds of lead as lead sulfate were obtained. The residue after lead extraction weighed 40 pounds and contained the gold and silver in addition to the SiO; content of the heads. Arsenic, of course, was extracted with the iron and the bismuth. was also in the final residue. In this particular operation no separate recovery of rare metals was made;

essentially ferrous sulfate with an admixture vof arsenic 1 and antimony and some slight amount of gallium, none of which had been noted in the initial partial assay of the ore. The residue at 82 was found to be substantially free of iron, contained no arsenic and only a trace of antimony. It also contained' traces of cadmium, bismuth and germanium as well as gold and silver, as found in the original partial assay.

The solid residue after the iron extraction was then passed to a flotation station as indicated at 83, the solids dispersed in aqueous meda in accordance with flotation practice, a relatively small proportion of creosote was added and the product rendered slightly alkaline by the addition of soda ash, and dotation carried out using a standard Denver equipment cell. A lead concentrate was produced as shown at 84, and this concentrate was found to contain 82% leadas the sulfide together with gold and silver, copper, and bismuth, of course, in small proportions.A The residue after 'the lead flotation was a high grade zinc concentrate as shown at 8 6 containing 61% zinc with traces of cadmium and germanium.

In a modification I passedthe roasted concentrate to a flotation station 87 (see the left hand side of FIG. 3 and the broken lines indicatingv modification) without preliminarily leaching the iron. I first produced a lead concentrate in the same manner as previously described using creosote -as a .flotation reagent and rendering the material slightly alkaline 'as previously described. By these means I produced a high grade lead` concentrate at 89 containing 70% lead. The concentrate, of course, had the usual small proportions of silver, gold, copper and bismuth, as previously described.

The aqueous residue after the flotation of the zinc concentrate was then treated with sulfuric acid to render the product slightly acid and flotation was again carried out using pine voill as a dotation reagent. This dotation step left the iron behind and produced a high grade zinc concentrate as indicated at 91 containing 54% zinc. The

tailings indicated at 92 contained substantially all of the iron as FeO together with arsenic and traces of antimony and gallium.

The results obtained in accordance with this Example employing expensive reagents, particularly collectors, to

producea commercial lead concentrate and alcomm'ercial zinc concentrate, both of which are sold asendprodu'cts of the mining-company to commercial smelters. The lead concentrate produced in accordance with these commercial practices will average about 30% lead, while YI obtain an 82% lead concentrate in one case and a 70% concentrate in the4 other. In the ca'se of thezinc concentrate, the commercial product will run about v50% zinc concentrate, while I obtain a 61% zinc concentrate in one case and a 54% zinc concentrate in the other case. The percentage of lead and zinc in thel concentrates does not tell the whole story,however, because the lead concentrate will contain up to 20% iron whereas my lead concentrate is substantially free. of iron', and the lead concentrate will also contain up to `to 12% zinc whereas my lead concentrate is substantially free of zinc.

may .run as high as 80% copper andusually it will have som'e` metallic copper associated with it. The tailings are an iron. concentrate as indicated. at .114 containing substantially all of the liron in the heads `and -being as high 4as 95% in iron as FcO making it an excellent furnace feed in conventional steel making operations. The iron oxide (FeO) of this concentrate may be treated in any one of many additional methods as already described hereinabove, but i-n any case it is afhighly valuable commercial product. The Icopper vsulfide may be treated in accordance with many diiferent procedures such, for example,

as delivery to a copper smelter for improving the character of the matte normally vfed to the converters. This Example 5 as illustrated in FIG.` 4 v-may be compared with practices of the prior art.- While it is known.

that copper can be extracted from simple complex suliides such as chalcopyrite, the ore or concentrate must first The commercial zinc concentrate is `somewhat better in that the iron content is only about 8% or 9%, as contrasted with approximately zero with the concentrate of my invention, but it will contain about one-half percent of lead and some copper, while my concentrates are substanf tially free of these two metals. thing is that in the commercialoperation discussed, the concentrates will only represent apart of thelead and zinc values'in the ore, whereas in-the case ofmy invention substantially 100% of the leady and zinc values in the ore will end up in the concentrate. One of the significant things about Example 4 is that it shows clearly that the complex is broken and a very clean separation between the various metals in the-heads is obtained.

Example 5 In accordance with another example (looking at FIG. 4) I selected an iron-copper sulfide from the Bagdad area of Arizona comprising principally chalcopyrite with some Still another important slight admixture of. products representing degradation of chalcopy'rite.v The concentrate. contained about 22% iron, 26% copper', 0.5% molybdenum and traces of gold, v

silver, selenium, gallium and germanium. The. ore was obtained as a concentrate. and whether or not itwas a complete bulk concentrate or whether or not lit was mined to take the ore as it came, I do vnot know, In any event,

the complex ironcopper`sulfide Acomprising.essentially chalcopyrite, as indicated at 92, was subjected to a low temperature. roast in accordance with the FIGQI procef dure as indicated'at93. The time of, roasting was 'threequarters of an hour, the maximum temperature 400 C.,

andthe proportion of cooled stack gases to air about six to one. vA Vsubstantial portion of the selenium was vaporized as indicated at 94 and recovered from the stack by a precipitation procedure,. The roasted product was then` subjectedl to a sulfuric acid Aleach as indicated at 96 -to remove iron in the form of a solution as indicated at 97, andthe residue washed at 98 and then passed into a higher temperature roasting zone at 99 where the remainder of the selenium was removed as shown at 101. The roasted product was then acid leached as at 102 and a'copper sul- `fate solution recovered as at 1 03, the residue 104 con'-` taining the gangue and the gold and silver present. The copper solution may be suitably purified as at 106 and a tie-polarizingv agent added asy required, after lwhich it is delivered to an electrolytic cell 107 and thel copper platedv out of solution onv suitable cathodes to produce electrolytic copperas indicated at 108.. The impurities removed from the copper may be processed to recovery as indicated at 109 and the electrolyte solution after a suitable pro-v, portion of lthe copper has been. stripped therefrom .as

shown at 111 is circulated back as indicated by the broken lines for furtheracid leaching at 102.

In a modification of the process as indicated by thel broken lines at the left of FIG. 4, the product resulting from the first roasting step is delivered to aotation station as indicated at 112,.andfa copper sulfide concentrate produced 'as indicated at 113. This Concentrate copper sulfide ore will run beween about a half percent and 'one percent copper, or in very good deposits up to about two percent copper, it is obvious that a substantial loss in the copper recovered cannot be tolerated. There is one thing to be said for pyrometallurgical methods as applied -to copper,- and that is that they recover a ve-ry substantial amount of the copper present in the concentrate,lusually in excess of 919.9% of such copper. This, however, is done at the expense of .recoveries of rare metals since, especially in the case of copper smeltin'g, the. very high temperatures employed in the vconverter will cause substantial losses. One reason for the smelting treatment i's that vat the very high temperatures involved the metal complex is broken, but the iron is 4still a wasted product, notwithstanding the fact that many attempts have been made to find ways of recovering it.

' In, almost all instances pyrom'etallurg-ical recovery of i copper is followed by electrolytic deposition, the principal reasons being that only in electrolytic cells are the gold and silver recovered, but also some other impurities in the final tough pitch copper produced at the smelter can b e removed by the electrolytic operation, and the pitch for further purification and poling `to remove the oxygen,

castingthe copper from the finishing furnace into anodes, and then shipping the anod'es to an electrolytic plant for the production of electrolytic copper.

, Thi's'operation may be compared with that of my present invention wherein a copper sulfate solution is produced directly from the concentrate and subst-antially one hundred percent'of the copper recovered. This copper solution after purification is then fed directly to an elecl o. trolytic cell without the need to interpose any of the expensive pyrometallurgical equipment now commonly employed. j y v l Looking at the left hand side of FIG. 4 I produce a 'very high grade copper sulfide concentrate, which makes an excellent -feed for a smelter, and at the same time recovers'ubst-antially all of the iron presentin the ore i-n the form of salable end products.

Examplev 6 FIG. 5 may be consulted in connectionwith thisExample 6. Starting with the iron-copper sulfide employed in the previous example as indicated at 116, the concentrate was subjected toa low temperature roast at 117 in accordance withV FIG. 1 using time, temperature and atmosphere conditions as described in connection with the previousA example. lFor convenience I have disregarded the presence of everything except the major constituents, except for gold and silver which in any case are found in the final residue. The roasted concentrate was leached at 118 to remove iron .in the form of a ferrous sulfate solution 119. The residue 121 contained the copper, and this was vsubjected to an oxidizing roast at a. temperature of 550 C. for one and one-quarter hours as indicated at 122. The resulting material was leached' as at 123 to remove copper in the form of a sulfate and produce a copper solution as indicated at 124. The residue after the copper removal at 128 contained gold and silver and represented from a percentage weight standpoint only a very small portion of the original heads. The ferrous sulfate solution 119 was treated to recover dry ferrous sulfate 121 which was treated in accordance with methods previously discussed to produce nely divided metallic iron -at 122. Some of this nely divided metallic iron was added to the copper solution at 124 and copper as the metal precipitated `at 128. This represents substantially one hundred percent of the copper in the heads. The copper, moreover, is highly pure and can be used directly without further purification except, ofl course, such compaction as may be indicated. The metallic iron added to the copper sulfate solution is converted to ferrous sulfate and this ferrous sulfate is withdrawn as a solution as indicated at 127. This solution, of course, is circulated back for further treatment in accordancewith whatever procedure has been established. In the tiow sheet all of the ferrous sulfate is indicated as being converted to Fe, lbut it should jbe borne in mind that only a relatively small percentage may be .so converted to produce a reagent for copper precipitation for step 124. It appears unnecessary to explain further in connection with this example the various ways in which the ferrous sulfate may be treated nor the manner of recovering the gold and silver and such other values as may be present in the residue of 126.

The process of the present example as indicated at FIG. may be compared with existing smelting procedures from both a recovery, cost and capital equipment standpoint. In the case of FIG. 5 the copper metal may be produced right at the mill Without the use of costly capital equipment such as those characteristic of smelting operations and electrolytic cells which, of course, 4are not only expensive, but must be located near available inexpensive power.v While certain of the features associated with this example are old, the procedure represents an over-all self-contained process in which the significant step is the controlled atmosphere roast, without which it would be impossible satisfactorily to accomplish the re` maining steps of the process.

Example 7 i I have already explained that I may conduct therelatively low temperature controlled atmosphere roast as illustrated iri FIG. 1 in such a manner as to solubilize more was thus produced as indicated at 136. In keeping with the initial guidelines I will not discuss the impurities in this solution, but it iso'bvious that in the normal ore there reference to other elements from this example, the mani ner of recovering such materials as cadmium, bismuth, and rare metals such as selenium, gold and silver having already been explained.

The concentrate indicated at 131 was passed continuously through a reaction zone in equipment of the type shown in FIG. l. The furnace atmosphere was controlled by delivering thereto approximately one part of air to three parts of cooled stack gases. The time required to pass the concentrate through the reaction zone was set for about fifty-tive minutes and the maximum temperature adjusted to about 600 C; The roasted product contained will be some trace materials which will nd their way into this solution and usually will have to be removed by stand- 'ard chemical means, although there are some impurities which can be delivered to an electrolytic cell without rais'- ing a problem of any kind. Depending upon the amount of iron present, this solution may be feddirectly to an electrolyticv cell, but preferably I remove some of the iron as ferrous sulfate, suitably by a crystallizing and precipitation process as indicated at 137. The solution is then delivered to an electrolytic cell 138 with at least somev of the original viron present as ferrous sulfate. This acts as a de-polarizing agent in accordance with known tech-y nology, and metallic copper is plated out on cathodes as l indicated at 139. The reclaimed solvent is returned for further use as a'leach as indicated by the broken lines.

Instead of depending entirely on control of the low temperature controlled atmosphere roast to solubilize all of the copper, I may solubilize only the iron or only a part of the copper and then entrain air by suitable means, such as bubbling the same beneath the surface of the solution tointroduce air with the leach solution and convert some of the ferrous iron to ferrie iron. The ferrie iron acts as a solvent for copper sulfide and suchtraces of copper metal as may be present. Copper sulfate is thus directly lcached by the acid and goes into solution with the ferrous sulfate. It should be pointed out in this connection, however, that if there is zinc or lead in the solution, this method involving the introduction of air should not be employed because it would have the elect of also solubilizing the lead and zinc sulfide and dissolving both of these materials into the solution as lead sulfate. -When producing an ironcopper sulfate solution in accordance with-this example I nd it convenient sometimes to partially solubilize the copper during the iow temperature controlled atmosphere roast and continue the solubilization by the technique involving oxidation of the' ferrous sulfate, I have found thatY all iron-copper sulfide ores do not behave in exactly the Vsame way. In some instances care must be taken to solubile only the iron, or otherwise some copper will dissolve with it, and in other cases only. the solubilization of the iron appears to occur readily, and very careful control is necessary to solubilize the copper as well. Generally speaking, however, despite peculiarities of the ore, it is quite possible on a `commercial basis to produce a solution from chalcopyrite comprising only iron sulfate or comprising bothI iron sulfate and copper sulfate.

Example 8 A concentrate of iron-copper sulfide ore comprising essentially chalcopyrite was provided at 146 as shown in FIG. 7. This ore was similar to that described in connection with FIG. 4, but it contained 23% iron, 27% copper and traces of gold, silver, selenium, gallium, germapercent, the time of passage through-the reaction zone was l about one hour, andthe maximum temperature was approximately 600 C. The resulting product 4contained relatively small proportions of FegOa, but most of the iron was in the form of-the ferrous oxide FeO. A very small perl centage of the copper was in the form of metallic copper,

3f the iron and copper were freely soluble in sulfuric acid.

Fhe roasted product was 'leached with sulfuric acid as' shown at v149, and for assurance of complete solubility, air was bubbled through the solution throughout the leaching itep and a solution was obtained as shown at 151 containng all of the copper asthe sulfate, substantially all of :he iron, plus the gallium and arsenic. The'arsenic was irst precipitated from the solution with mag'nesiurn, and metallic iron then added to the solution at 152 to precipitate .ubstantially one hundred percent of the copper. This :opper after washing was found to be a highly puremetalic copper suitable for commercial use as recovered., 'The errous sulfate solution as indicated at 154 contained subtantially all of the iron in the heads plus the ironintroluced to precipitate the copper, all this iron being present is' the ferrous.l sulfate. For all practical purposes the steps of the method hereofore described comprise a complete process in that all if the lcopper has been recovered as the metal and all of he iron has been recovered as a ferrous sulfate soluion which can merely be dried to produce a salable agri- :ultural grade ferrous sulfate having a ready sale, This art of the process can be comparedto smelting: procedures n which substantially onlythe copper is recovered and the ron discarded as a slag. The very definite advantage in he procedure disclosed, however, is that practically one' tundred percent vof the copper is recovered in accordance Vith very simple, inexpensive procedure requiring no highy expensive capital equipment and the iron normally a vaste product is recoverable asa valuable by-product.

My invention providesfor still additional recoveries, towever, which are not characteristic of copper ore. procssing exceptQof course, for the recovery of some ofthe :old and silver commonly picked up in the mud at'the `ottom of the electrolytic cell. Looking at station 156, he residue from the iron and copper sulfate leach conains the gold and silver and tracesof germanium and iolybdenum. The residue may be treated with a cyanide sgshown at,157,'gold.and silver recovered at 158, and

residue free of `gold and silver obtained as indicated at 59. This residue is then silver treated with a hot concenated sulfuric acid solution and germanium and molybenumrecovered at 161, germanium as the sulfate and iolybdenum as molybdic acid. As previously described, n-alkali metal hydroxide leach may also be employed at iis point. The residue at 162 comprises essentially only iO2 and such other gangue constituents as may be resent. v

The impurities in the ferrous sulfate solution-bothv as it ppears at 151 and 154 may be recovered in various ways. fhe ferroussulfa-te solution stripped of all of its copper s at 154 maybe delivered to an impurity recoverystation Preferably at least some elemental iron in the formv of very finely divided particles is produced for'use in the copper precipitation'step. Station 165 is indicated as the point of gallium removal, and the two-faced arrow between 165 and 166 indicates balanced iron and gallium removal as explained'.

In FIG. 7, for easy comparison with copper recovery y procedures of the prior art, I have shown in full lines the steps employed in the recovery of the copper which, as pointed out, is substantially the only product of any consequence recovered by conventional smeltng methods. All of the stations shownin broken lines represent-additional recoveries from a common type of iron-'copper sulfide, all of whichfare new with the present invention and represent recoveries not normally possible and, in many instances, completely impossible with conventional smelting methods. v

A bulk concentrate 171 (.FIG. 8) of a complex sulfide has been eliminated in this example as it is shown in FIG.

8. .This low temperature-roasting step solubilzedthe iron and the roasted ore was leached at 173 .to form a ferrous sul-fate solution at recovered aty 176.

The residue from the acid leaching to remove iron was again treated ata relatively-low temperature in the general mannervshown -in FIG. 1 and as illustrated at 177 in the drawing.- Conditionswere maintained approximately as Ain the tirst step at'172, but the furnacetemperature was controlled so that the maximum temperature immediately before discharge was about 550 C. The residue so heated was then treated `with concentrated. sodium carbonate solution at'178 to cause a double decomposition reaction involving the lead sulfate and thesodiumcar- 174 from which the iron was bonate, a sodium sulfatesolution being withdrawn as at v 179 and the residue passedto station 181l where it was washed and then leached withacetic acid to form alead acetate solution `'1 82, Lead was suitably recovered from this solution at 183, in this case -by adding sulfuric acid to precipitate lead sulfate and regenerate the acetic acid residue containing lead carbonate.

lch as 'at 163. By several dilferent means the initial :ach solution as at 151 can, if desired, be treated to reiove some of the impurties and, if this should be -th'e ase, such impurities may also be delivered tostation 63. .The solution then atv163 .will comprise the ferrous llfatesolution from station 154 from which all of the Spper has been stripped, and the impurities willy 4comrise essentiallygallium. The gallium may be removed y several different procedures. A very satisfactory lethod is to crystallizel substantial portions of ferrous tide from solution to produce a solution richer in 'galum. This solution is re-circulated to build up the galum vto a relatively high figure, and the gallium then prepitated with ammonium acetate. A basic gallium ace- ,te is precipitated (with some iron), the product further lried, and the basic lgallium acetate heated to produce galliuin oxide which may then be treated by known ,eans to produce a final product. The 'solution with me o r all of the impurities removed at l164 is shown i being delivered to an iron recovery station-166 whereY on may be recovered in accordance with any suitable, asirable, salable form such as discussed hereinabove.v

which was then' available for further leaching of additional It is, of course, obvious that many different ways of recovering the lead from the lead acetate are lavailable and the lead may be treated to'meet anyv commercial demand such, for example, as

The residue from-the acetic acid leach at 181 was then washed and dried and again roasted as at 184 in a controlled furnace atmosphere using equipment as shown at FIG.1, andthe same general conditions as employed in.

the first two stepsv indicated at 172 and 177, except that the maximum temperature was raised to about 600 C. The s o-roasted oreafter cooling was leached with dilute sulfuric acid as shown at 186 to produce a copper sulfate solution as shown at 187. The copper solution was delivered to a copper recovery area 188 where some of it was -treated'in an electrolytic cell as shown at 1 89 to produce electrolyticcopper, and some delivered to a tank at 191 where zinc was added to precipitate copper as -themetal and form a solution ofzinc sulfate.

The lresidue at 192 was washed and dried and then subjected to a controlled atmosphere roast at 193'using the same general conditions as described for the previous roasts, and the samel kind ofequipment but with the 25 maximum temperature controlled to about 650 C. This roast was then leached with sulfuric acid at 194 to produce a zinc sulfate solution 196, delivered to a `zinc recovery station 197 where the zinc was recovered in accord` ance with several dilerent procedures. l

In addition to the metallic zinc which is readily prepared from this solution there are several zinc compounds having ready commercial sale, such as zinc sulfate itself and zinc oxide. Zinc oxide is not only used in medicine, but in many areas it has extensive use as a white pigment.

The residue from the acid leach to remove zinc is shown at 198. v as shown at 199 to produce a gold and silver cyanide solution at 201 which is then readily treated in accordance with known methods torecover gold and silver at 202. The residue at 203 represents only a very small percentage of the original heads and comprises substantially only gangue material, except in such instances Where therevare traces of rare metals such as germanium to be found in it.

Example 10 A complex ore comprising essentially iron-zinc-lead sulde with a relatively small proportion of copper, about 0.7 ounce of gold per ton of concentrate, and about 24 ounces of silver per ton of concentrate was treated by means of a conventional high temperature roast. The

metals .were then taken into solution and only 60% of the zinc was recovered and 58% of the lead. The solution also contained only a trace of copper. The residue after acid leach was treated with a cyanide solution for ten hours, but only approximately 40% of the gold and silver was recovered. s

This same ore treated in accordance with my invention and specifically in accordance with the general ow sheet shown in FIG. 2 accomplished, substantially complete recovery of the zinc, lead and copper, and slightly more gold and silver than the assay showed tobe present.

Example Il A commercial lead concentrate was selected which assayed:

23% iron 32.9% lead 1.9% copper 9.4% zinc 0.71poz. gold per ton 34 oz. silver It was also found that it contained 1.6% arsenic. This lead concentrate which represented a final product of the prior art was then heated at 450 C. in a controlled atmosphere in the manner shown in FIG. l, the propor tion of cooled stack gases to air delivered to the furnace being about ve to one. After roasting the iron was ex-` tracted with dilute sulfuric acid, the material being subjected to a simple grinding operation during the leaching procedure. Substantially all of the iron was removed.

The residue after removal of iron was then subjected It may be treated with a cyanide .solution casami- 2d Example 12 A concentrate of Va chalcopyrite ore from the Bagdad district of Arizona contained:

vPercent Iron s 29 Copperl c 24 Sulfur 30 4Molybdenum 0.8

and a balance of calcium carbonate with traces of other impurities. This concentrate was roasted in accordance copper leached with sulfuric acid, yielding a very pure copper sulfate solution. The copper was precipitated as metallic copper with finely divided metallic iron.

Example 13 Using the same concentrate as in the preceding example, the roasting time was increased to one and onehalf hours, the temperature increased to 500 C. and the proportion of stack gases to air cut in half. On leaching with sulfuric acid substantially all of the iron and copper were taken' into solution in the form of sulfates. This solution was treated in two ways, in one of which the copper was precipitated yout of solution as metallic copper by means of metallic iron, and in the other of .which the solution was delivered directly to an electrolytic cellv and a substantial proportion of the copper stripped from the solution and plated on a cathode, the ferrous sulfate acting asA a depolarizing agent.

For those skilled in the art to have a full understanding of the importance and scope of the present invention, further consideration should be given toA the rare metal content of complex ores and the recovery thereof.

Modern technology continues to develop new uses. for rare metals such as gallium, indium, rubidium, germanium, selenium, tellurium, thallium and the like. To a considerable extent commercialization is frequently delayed or even prevented by shortage of supplies. At the same 'time manycomplex ores containing such metals are processed without recovering such metals, and many .ores containing substantial amounts of such metals are not even capable of economic beneficiation. Even such borderline metals as antimony.. bismuth and cadmium are frequently in short supply,.while fairly substantial proportions of such metals will be found in tailings and on slag dumps.

o My present invention makes possi-ble substantially complete recovery', frequently on `almost a lay-product cost basis, of all rare metals of any kind found in the ore. It is frequently surprising how many different rare metals willbe found in a single ore, with no single one of such f rare metals included in standard assays.

Asa very cogent example, I Wish to refer to the ore described in Example 2. From one ton of concentrate assaying as shown on the top of page 30, the amount of rare and borderline metals recovered was as follows:

Gallium 6 pounds (as oxide).

Iridium 3.5 pounds (as oxide). Germanium 846 grams (as oxide). Selenium 2 pounds-(as metal).

Bismuth` 1.2 pounds (as metal). Cadmium 1.8 pounds (as metal). Antimony 17 pounds (as metal).

neral containing the rare metal, but they appear obiusly, in at least many instances, to be so bound chemlly and/or physically that recoveries have been for the st part impossible by previous methods. vNor amA :I are of the exact effect of the processing by my present ention on such rare metals. Whatever the mechanism, y are released so that they become freely available at ne step of the process. I have shown, or explained, tthe base metals after the first roast will be FeO and sulfides of the other base metals. Iam not aware at -bination of complex chemical and low concentration tends to make assays low and recovery still lowers My process breaks the complex and passes'on all of the -precious vmetals (without overheating) to a final concentrated residue in which the percentage of precious metals is in- `creased at lleast forty-fold.

, As an example, there is a property in northern Arizona which had been worked to recover the oxide ore for gold and silver recovery, and .the sulfide portion thrown on a dump because of inability to extract gold and silver -on a commercial basis.- I made a simple bulk concentrate of this-dump material, 4the concentration being about five present time exactly what loccurs'in the case of the e metals, but a very large number of experiments ws that they are always substantially one hundred cent recoverable. There is a significant point as to recovery of germanium, for example. Assuming that re is 0.005% germanium in the original ore, in the il residue from which germanium may be extracted, re has been a concentration of at least forty times, and germanium concentration will thenbe over 0.2%;v It Lhen freely leachable, for example, with sodium hy- `xide solution and may then be freely removed from ation by hydrolysis.

t least a substantial part of the indium is lfound in the c sulfate solution. Jtion, there would be a co-precipitation of iron and ium and recovery would be very different. 1 present, as in the zinc sulfate solution ofthe present ention', the indium can be recovered as a pure product. the zinc sulfatev solution, there is normally-a fair aunt of `both indium and cadmium. By adding zinc Lal, indium and cadmium will precipitate together as ials. They are then again dissolved in dilute sulfuric i and the solution neutralized till it shows just al- ,ne to methyl red. 'After standing-for several hours,v

If there were any iron in. this zinc With no by warming,"indium separates out as the hydroxide t this can be C. Vhat isjsaid` of-indium is true in part of gallium. It normally be found also in the zinc vsulfate solution, is not precepitated with indium and cadmium. To

love galliu-m the solution is hea-ted after removal of' .um vand cadmium, and then made slightly alkalinel by itiori of zinc oxide. .The solution is then cooled and ium precipitates out as zinc gallate (ZnGa2O4). Gali may be precipitated as the sulfide-(Ga2S3) by dis-l 'ing the zinc g'allate in sulfuric acidleaving about two :ent free sulfuric acid, and the gallium sulfide may :alcined or treated in several d ifierent ways to produce al or other desired compounds.

reviously I noted that in practically all cases er, than shown by assay of the head materialto be tent. This may be contrasted' with other processes which almost'invariably :recoveries will be less than wn by assay to be present. In some cases this'differa is marked and there are instances, particularly in case of silver, where recovery of precious metals wn to be presentis for all practical purposes 'imposhere appears to be at least two reasons for explaining above observed facts. Apparently, and this is one aced theory, silver particularly will be locked in a comof the type called sulfo minerals to distinguish n usual sulfides. They are characterized by the formu- YZS,t, in which X, Y and Zl are cations, one of h is silver, S is sulfur andx is a variable. Precious als are present in very small percentage, and the comreduced to the met-al by H2 at aboutv to one. This assayed about 415% iron, 33% zinc, 13% copper and 2% lead, and sixty dollars per ton of gold and silver. By my process the base metals were readily vrecovered, and I also was able to recover all of the gold and silver present.

In the practice 'of my invention, due recognition may be given to practices in the industry.v As an example, the United States Mint frequently prefers to recover gold and silver in the form of high concentrates, rather than in the form of bullion or ingots, which must be re-processed anyway. 4ln detailing the stepsof Imy proces's,.I

refer to removal of gold and silver by cyaniding or other treatment of the final residue. -It must be remembered, however,` that the concentration is such that such residue may' be shipped direct to the mint and full valuereceived for its gold and silver content.

Early in the specification I referred to the iron after the first roasting step' as being acid-soluble. In most of the examples, however, I have referred to a sulfuric` acid leach because use of sulfuric acid is common in leachingprocedures and, moreover, my process provides a ready'source of sulfuric acid. It should be remembered, however, that the iron is leachable 'by means of most inorganic acids and a relatively very large number of organic acids, including most ofthe common ones. Also, in various steps of the process the` examples are illustrative and not limiting.` Illustratively, wherereference is madev to the use of sodium carbonate to produce lead carbonate,

other comparable carbonates, such as potassium carbonate, may of course be used. v

Those skilled in the art willunderstand that because of the breaking of the sulfide complex in the first controlled atmosphere relatively low temperature roast, and

` because of the relatively large number o f deposits of complex o res, no two of which as a rule are absolutely alike, there are large numbers of ow sheets which may be engineered to process such lores without departing from the scope of m'y invention. Illustratively, according., to

. some specific"examples,-1 produced concentrates of iron,

Il 0b# ed greater recoveries'of precious metals, particularly leadfand zinc, each with certain inclusions, and all produced by means of the .simplest kinds .o fiiotation techniques. It-should be borne in mind that by means of the most advanced and latest't'echniques involving'new frothers and sensitive selective collectors,` I may produce still higher grade r'concentrates than some of those indicatedeven concentrates comprising essentially only the metal desired in the form of its sulfide or other compound. Itmust be remembered that by means-of my process the sulfide complex, which heretofore has prevented complete separation of metal values, is broken, and I may effect then almost perfect separations in many instances.

Illustration ofA a marked departure from the illustrative flow sheets shown in the drawings and the examples given in the specification may be found inthe recovery of precious orrare metals. In the examples given the prime purpose is the recovery of base metals such as iron, copper, zinc and'lead.. There may 'be many instances, however, such, for example, as the low content 'of base metal or location of the ore, where processing and recoveries .are directed principally or even only to precious and/or rare metals. In such a casethernovelty ofthe present invention would be directed primarily to such purpose and the ow sheet modified accordingly. Many changes may be made if precious metal and rare metal recovery are not approached on a by-product basis. Thus, by controlling the iron leach with dilute sulfuric acid after the controlled atmosphere roast to a pH of between l and 2 at the end of the leach, I may remove a great deal of the gallium and indium as sulfates. Indium can be precipitated from solution with finely divided iron', and gallium as a basic acetate or ammonium compound. It may also be precipitated as the sulfide with hydrogen sultide (H28). Germanium, which normally is found kin the residue after the final leach, may be taken into solution at an earlier point in the cycle and also precipitated as the sulfide.

Terms in the specification and in the claims are, in general, used in their usual sense in the art unless the 6. A process of treating a complex sulphide ore con- Y taining iron which comprises: v

context indicates to the contrary. Thus, when I refer to iron and other base metals in the ore, I do not imply the presence of a chemical base such as sodium hydroxide, but a common metal as distinguished froma rare metal. When I refer to processing an ore, I include not only the ore itself but any fraction or concentrate thereof, even though l prefer to produce a bulk concentrate Aof the ore 'as the initial step in the practice of my invention.

Finally, although I have explained and illustrated my invention in detail so that those skilled in theA art will understand the manner of practising the same, the scope of the invention is defined by the claims.

I claim':

1, A process of treating a complex sulphide ore containing iron to obtain the metal values therein which comprises:

(a) heating said ore at a `temperature between 400 and (b) in an 'atmosphere comprising a minor portion of air and a major portion of sulphur dioxide,

(c) until substantially all of the iron present is in the form offerrous oxide,

(d) and cooling the roasted productvont of contact with air to a temperature below 300 C.,

2. A process of treating a complex sulphide ore containing i-ron to obtain the metal values therein which com` prises:

(a) heating said ore at a temperature between 400 and (b) in an atmosphere comprising a minor portion of air and a major portion of cooled stack gases, consisting essentially of a mixture of sulphur dioxide and nitrogen,

` (c) until substantially all of the iron present is in the form of ferrous oxide,

(a) continuously passing said ore through a sealed heated reaction zone, f

(b) continuously discharging roasted ore at a discharge end of said reaction zone,

(c) passing a relatively small proportion vof air into the reaction zone to partially oxidize the ore as itmoves through the reaction zone,

(d). continuously removing hot reaction chamber gases from the reaction zone at the ore receiving end thereof, Y

(e) delivering a major amount of sulphur dioxide and nitrogen-to the reaction zone at said ore discharge end, whereby to dilute said air and produce a furnace atmosphere high in sulphur dioxide, and

(f) controlling the temperature of the ore to cause a gradual rise thereof to a minimum temperatureof 'about 400 C. and a maximum -temperature of (g) controlling the rate of movement through the reaction 'zone to convert at least substantially all of the' iron present to ferrous oxide and,

(h) cooling the roasted product out of contact with air to a temperature below 300 C.,

(i) and acid leaching the resulting roasted product to recover substantially all of the iron;

7. A process of treating a complex sulphide ore of a class including iron, lead, copper and zinc which comprises:

(a) heating said ore ata temperature between 400 and 600 C., (b) in anV atmosphere comprising a minor portion of air and a major portion of sulphur dioxide,

(c) until substantially all of the iron present is in the `forrn of ferrous oxide,

.'(d) and cooling the-roasted product out of Contact with air to a temperature below 300 C.

8. A process of treating a complex sulphide ore coritaining iron, copper, lead and zinc to obtain the metal (d) and cooling the roasted product out of contact with air to a temperature below 300C.

3. A process of treating a complex sulphide ore containing iron to obtain the metal values therein which comprises: v

(a) heating said ore at a temperature between 400 and (b) in an atmosphere comprising a minor portion of 'air and a major` portion of sulphur dioxide,

(c) until substantially all of the iron present is in the form of ferrous oxide,

(d) cooling the roasted product out of contact with air `to a temperature below 300 C. and,

(e) acid leaching the resulting roasted product to recover substantially all of the iron.

4. A process of treating a complex sulphide ore containing iron to obtain the metal values therein which comprises:

(a) heating said ore at a temperature between 400 and (b) in an atmosphere comprising a minor portion of air and a major portion of sulphur dioxide,

(c) until substantially all of the iron present is in the form of ferrous oxide,

values therein which comprises:

`(a) heating said ore at a temperature between 400 and (b) in an atmosphere comprising a minor portion of I air and a major portion of sulphur dioxide,

(c) until substantially all of the ironV present is converted to ferrous oxide while substantially all of the other metal valuesin the ore remain as sulphides,

(d) andcooling the roasted product out of contact with 'air to a temperature below 300 C. v

References Cited bythe Examiner BENJAMIN HENKIN, Primary Examiner. 

1. A PROCESS OF TREATING A COMPLEX SULPHIDE ORE CONTAINING IRON TO OBTAIN THE METAL VALUES THEREIN WHICH COMPRISES: (A) HEATING SAID ORE AT A TEMPERATURE BETWEEN 400 AND 600*C., (B) IN AN ATMOSPHERE COMPRISING A MINOR PORTION OF AIR AND A MAJOR PORTION OF SULPHUR DIOXIDE, (C) UNTIL SUBSTANTIALLY ALL OF THE IRON PRESENT IS IN THE FORM OF FERROUS OXIDE, (D) AND COOLING THE ROASTED PRODUCT OUT OF CONTACT WITH AIR TO A TEMPERATURE BELOW 300*C. 